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By R. H. Nafziger

Bureau of Mines research has shown that very low-sulfur pressure vessel steels with improved mechanical properties can be prepared by two melting processes. Four pressure vessel steels having sulfur contents ranging from 0.0008 to 0.0040 wt-pct, were produced by the electric-arc furnace or by the electroslag processes. Calcium carbide was used as a desulfurizer in the electricarc furnace. Fluxes containing at least 40 wt-pct CaO provided sufficient desulfurization during electroslag remelting. Element recoveries were higher for electroslag remelting than for metal melted in the electric-arc furnace. Bar rolled from electroslag-remelted ingots generally possessed higher tensile strengths and ductility than bar rolled from it t melted in the electric-arc furnace. Charpy V-Notch upper shelf energies were also improved as a result of electroslag remelting. The primary failure mode during impact testing was was microvoid coalescence. Sulfide inclusions were minimal, especially in the electroslag-remelted material.

Jan 1, 1979

By R C. A Minnitt

The two methods for calibrating the parameters K and (alpha) for use in Gy’s equation for the fundamental sampling error (FSE), the duplicate sampling analysis (DSA) and the segregation fractionalised analysis (SFA) are described in detail. A case study using identical broken reef material from a Witwatersrand-type orebody is calibrated using the DSA and SFA methods and the results are compared. Classically, the form of Gy’s equation for the FSE raises the nominal size of fragments given by dN to a power of 3. Gy’s formula is a theoretical formula applicable in the case when fragments are randomly collected one by one, which is obviously never the case. This is because the fragment shape factor (f) and the granulometric factor (g) were introduced to take care of this anomaly. More recent formulae, such as those suggested by François-Bongarçon and Lyman, are empirically based formula. A later modification of Gy’s formula raises dN to the power ? (alpha), the latter term being calibrated with the coefficient K in the DSA and SFA methods. The preferred value of ? (alpha) for low-grade gold ores used by sampling practitioners in the mining industry is 1.5. A review of calibration experiments for low-grade gold ores using the DSA and SFA methods has produced values of K that vary between 70 and 170 and values of alpha in the range 0.97 to 1.30. The average value for alpha is found to be 1, rather than 3 as originally proposed in the classic form of Gy’s equation or the industry preference for 1.5. It is suggested that for low-grade gold bearing ores, the equation for the FSE should raise dN to a power of 1. Such an equation for the variance of the combined FSE plus the grouping and segregation error, now referred to as the short-term component of the quality fluctuation error (QFE1), greatly simplifies the characterisation of gold ores, which now only requires the calibration of K for a given mass and established fragment size. The implications for the heterogeneity test from the generalised equation are that provided the fragments have been screened to within a narrow size range, any particular size will return a value for K that is acceptable for use in the sampling nomogram.CITATION:Minnitt, R C A, 2017. A generalised form of Gy’s equation for gold ores – empirical evidence, in Proceedings Eighth World Conference on Sampling and Blending , pp 331–350 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By J. R. Soule

"The fluorspar deposits in the Gallinas Mining District, Lincoln County, New Mexico, were examined early in February and again in April 1943 by the Bureau of Mines because of the possible importance of newly discovered deposits and of old mine workings showing fluorite-bearing material. There &re many of these deposits spread over a considerable area, and it appeared that exploration might indicate sufficient reserves to support a local mill that could contribute to the supply of fluorspar needed for war purposes. Accordingly, the Bureau initiated an exploratory program in June 1943, which was completed in August 1944.H. E. Rothrock, Roy Earhart, R. G. Smalley, and V. C. Kelly, geologists of the Federal Geological Survey, began mapping and studying the district in May 1943 and continued this work throughout the Bureau's exploration.David J. Stauffer, district ranger, U. S. Forest Service, and several claim owners assisted in the examinations and subsequent exploration.The fluorspar deposits in the Gallinas Mountains are reached by an unpaved forest road, which, leaves U. S. Highway 54 half a mile south of Gallinas station. The road runs due west for 3 miles, then turns due north just beyond the first cattleguard. This north branch is followed for 2 miles to a second cattleguard. There a left branch extends northwest up Red Cloud Canyon to the Red Cloud, and. All. American deposits; the right branch continues north for 2.miles past the Wade Porter ranch, to another left branch, which goes almost 2 miles west to the Hoosier Group deposits. The remaining deposits are reached by trail from either the Hoosier Group or, more easily, from the Red Cloud claim (fig. 1).The Hoosier Group is in sec. l9,, T.. 1S.,1 R. 12 E.; all of the other known ore occurrences are in secs. 13, 14, and 23 to 27, inclusive, T. IS., R. 11 E. All of the deposits are in the Lincoln National Forest and within the limits of a State game refuge. The nearest railroad loading point is Gallinas, Station on the Southern Pacific Railroad. Corona, the next station north of Gallinas, is a Mile or two farther from the deposits.Minor supplied, telegraph, telephone, freight, and express service, and limited living accommodations, Are, obtainable at Corona, 10 to 15 miles from the deposits of Carrizozo, N. Mex., 50 miles south."

Apr 1, 1946

By J. M. Singer

This study was undertaken to obtain some fundamental information about hazardous mixtures of coal dust and methane (firedamp) likely to occur in mines. Fuel concentration limits for ignition of mixtures of coal dust, methane, and air have been determined by a hot-gas ignition method. The limiting concentration of each of the two fuels was usually less than that corresponding to its lean flammability limit in air. The lean limits of flammability of coal dust vary with the type and size of the dust; in this study coal dust concentrations were less than 80 mg/liter. Ignition jets were turbulent pulses (emerging from a channel of 0.5 cm diam) resulting from explosions of stoichiometric mixtures of methane, oxygen, and nitrogen, Temperatures of the ignition jets were varied by changing the oxygen index (OI) = O2/(02 + Nj2) of the primary methane-oxygen-nitrogen mixture (adiabatic flame temperatures from 2,000° to 3,000° K). The objectives of the study were to establish quantitative relations between concentrations of each fuel at ignition limits of hybrid mixtures and to study the mechanism of ignition of these hybrid mixtures by hot turbulent gases, Empirical equations were fitted to concentrations for four coals, namely) Pittsburgh Seam (sizes A, B, and C), Sewell No.2 Seam, pocahontas No.3 Seam, and anthracite. Equivalences of coal dust to methane were computed from these equations; equivalence being defined as -Ac/Am, where c is the coal dust concentration and m is the methane concentration, each in milligrams per liter of air. It was found that the equivalence increases as the temperature of the hot-gas jet decreases. In general, the equivalence is also an inverse function of volatile content of the coal and a direct function of the particle size, However, volatile con-tent in itself is not a sufficient correlating variable, nor is the specific surface area of the coal dust.

Jan 1, 1967

By E. A. Clarke

The vapor-phase section of the Coal Hydrogenation Demonstration Plant at Louisiana, Mo., was operated for the first time in April 1949. The run lasted only 10 days because of time limitation (pending plant dedication in May) and the small amount of charging stock available. The primary purpose of the operation was to break in and determine the operability of the equipment and to supplement the operators' classroom training with actual practice by operating the plant under mild conditions. A secondary aim was to obtain data on the hydrogenation of petroleum oil and lignite tar distillate at approximately 10,000 p.s.i. while making gasoline and Diesel fuel oil. Although the operations were carried out under conditions considerably below those used in the plant design, a great deal of experience and data was obtained. As it is not feasible at this time to conduct further vapor-phase operations on similar stocks, this report has been prepared to present and summarize the data obtained through this operation.

Jan 1, 1950

By W. I. Nelson

This paper describing the mining practices at the Engels mine is one of a series being prepared by the United States Bureau of Mines on mining practices, methods, and costs in the various mining districts of the United States. The Engels Copper Mining Co. operates two mines about 3 miles apart in Plumas County. Calif. Both mines send ore to the same mill, and the methods of the larger, or Engels mine, are described in this paper. Between 800 and 1,000 tons of 1-3/4 to 2-1/2 per cent copper ore are mined daily from this mine by a shrinkage method. Connection is made with the Western Pacific Railroad at Paxton by the Indian Valley Railroad, which is 22 miles long and operates a daily mixed train to the mine.

Jan 1, 1930

By John H. Soule

The Torpedo Copper mine in the Organ Mining District, Dona Ana County, N. Mex., was discovered In 1899. Production was begun immediately and continued until 1907. Since 1907 very little ore has been produced, although Many individuals and companies have been interested in the property at one time or another. Approximately 6,621 feet of diamond-core had been completed on the property before 1949. The results from some of this drilling were unsatisfactory because of the poor core recovery. In addition to its investigation of the Torpedo ore deposits, attention was given by the Bureau of Mines to obtaining better core recovery through the use of special equipment and techniques. This project was begun October 3, 1949, and was completed April 30, 1950. Four diamond-drill holes totaling 2,140.5 feet, which includes 18.2 feet of overlap in a deflected hole, were drilled. All drilling was done with a Government-owned hydraulic-feed drill with some special rods and core barrels, in addition to the usual equipment. The over-all core recovery was 89.5 percent. Omitting the drilling in overburden where little or no core was desired, the recovery was 94.6 percent. Core recovery in the ore zones averaged 95.2 percent. This paper describes the location, the general features of the area, the geology, the Bureau of Mines procedures, and the results attained by these procedures.

Jan 1, 1952

The intent of this manual is to provide guidelines for inspectors of drilled shaft construction otherwise referred to as drilled piers, caissons or bored piles. It should also be helpful to train beginning geotechnical engineers with limited first hand experience or for geotechnical engineers who are training inspectors. It is expected that the general guidelines and procedures described here will be supplemented by local geotechnical engineers working in a particular region based on local practice and experience. Since drilled shaft construction procedures vary widely based on local geology and size of project, an all-inclusive manual is not considered practical. This manual focuses on traditional topside inspection for routine drilled shaft construction. Other inspection and evaluation methods that rely on electronic equipment are available, but these methods require extensive training and experience in the interpretation of results. These more sophisticated methods of drilled shaft inspection and evaluation are common in construction, but they are usually provided by specialty subconsultants. The subconsultant will typically submit a report for each shaft, which will be reviewed by the geotechnical engineer. Other manuals are available to describe the electronic methods of drilled shaft inspection.

Jan 1, 2004

By W. D. Musgrove, B. W. Gandrud, A. C. Richardson

"INTRODUCTION The Cahaba coal field is the second largest coal—producing area in Alabama. This field underlies parts of Shelby, Jefferson, St. Clair, and Bibb Counties and has an area of about 270 square miles. The Cahaba formations contain a number of coal beds, about 8 of which are mined at present. Although the area contains many folds and faults, samples from various mines and beds indicate that, except for the variation in ash content, the coal from this field is very uniform in character and quality. It is a high-volatile bituminous coal used chiefly for steam-raising purposes and domestic fuel. The coal as mined usually contains too high a percentage of impurities to be marketed without some preparation. The majority of the mines have washing plants in which jigs are the favorite type of equipment, although at some mines tables are used for cleaning the slack sizes. The washing characteristics of the coal from the Cahaba field are very uniform and differ from those of the near-by Warrior field, in that the coarser sizes contain a much lower percentage of bony material than the finer sizes. In general, the bone increases with decrease in size. It breaks into thin flakes, making difficult its elimination in the cleaning process, and gives rise to what is known among the washery operators of the Cahaba field as the ""rash problem.""A previous publication 5 referred to rash from the standpoint of jig operation and suggested some means of handling it with jigs. This report presents the results of table tests on two samples of fine coal from the Thompson and Woodstock beds, respectively, of the Cahaba field to determine the influence of rash on table efficiency.This investigation has been made in connection with the study of coal-washing problems now being conducted by the U.S. Bureau of Mines and the University of Alabama, with the cooperation of the different coal operators of Alabama."

May 1, 1934

By Perry D. Teitelbaum

In two earlier studies the Federal Bureau of Mines undertook to explore comprehensively the dimensions of the Nation's energy balance. The first of these studies, published in 1950 as Bureau of Mines Information Circular 7582, by Harold J, Barnett, was primarily concerned with projections of future domestic energy requirements and supplies, because these related to projected growth in major fuel-consuming sectors of the economy, changing rates of energy-use efficiency, and interfuel substitution patterns. But, to provide the essential background for these projections, it was first necessary to assemble and examine in some detail the quantitative information on the size and shape of the current United States fuel economy. Among other things, Barnett's path-breaking study indicated the need to consider energy as an entity, particularly for public and private policy makers, industry and commodity analysts, economists, and others concerned with long range questions of domestic and international energy supply and demand. The study also spotlighted a gap in the fabric of current economic intelligence; no comprehensive collation and reconcilation in any single report of available published statistical information covering the origin -a disposition of all energy materials existed. In response to this latter need, the Bureau of Mines prepared the second of the two studies referred t3 above. This was published in 1951 as Report of Investigations 4805, by W, H, Lyon and D, S. Colby, and it consisted of statistical information on the production and use of each of the various fuels and electric energy in the United States for 1929, 1939, and 1947; all were presented in tabular form and on a uniform basis. The three years selected were chosen largely on the basis of data availability.

Jan 1, 1961

By R. R. Lowery

An analysis of three 1-1/4-diam manufactured wire ropes of 6x19 classification was done to identify and describe reference conditions of new rope. Reference condition is of value when examining tested or retired ropes. Measurements of strand and rope pitch and of wire and rope diameter confirmed adherence of all ropes to those construction specifications. Mechanical properties of strength and ductility were determined by accepted methods for the ropes and wires from the ropes. Individual wires from one rope failed to meet minimum wire ductility values. Wire chemical composition varied in average carbon levels, from 0.70 to 0.78 wt pct, and in manganese content, from 0.45 to 0.80 wt pct. One manufacturer's product contained 1 to 3 wt pct Ni and Cr. Heavily wrought pearlitic steel microstructures were porous and contained inclusions. Cold work variations were evidenced by wrought grain width. Recorded anomalies reflected the condition of the manu-factured products. The report provides data that compare the three wire ropes studied, but cannot verify the effects that these conditions or anomalies have on rope service life.

Jan 1, 1983

By Arnold B. Hubbard

A thermal decomposition study of the kerogen in three Colorado oil shales was made at different temperatures by heating the shales in the absence of oxygen at atmospheric pressure and measuring the decomposition products (gas, oil, and bitumen) formed. From these data the specific reaction rates, "k", for the thermal decomposition of kerogen were determined for each temperature studied, making it possible to calculate the time required to accomplish a given amount of decomposition within the temperature range. The influence of temperature on reaction rates was presented by plotting the data according to the Arrhenius equation, thus providing a means of estimating the values of "k" for any temperature within the range of this study. It was found that the specific reaction rates for the thermal decomposition of the kerogen in the three sampler were comparatively equal at corresponding temperatures.

Jan 1, 1950

By Wallace W. Roepke

The Bureau of Mines tested new designs of drag bit cutters with sintered diamond inserts instead of standard tungsten carbide tips. The bits were tested for orthogonal cutting forces, primary dust generation, and incendivity, with a standard plumb-bob type conical cutter (600 included tip angle) as the reference. Preliminary wear and impact-failure testing were also done. The results were mixed. Two new designs had forces similar to those of the reference bit, but one new design had about twice the normal force. This same design had almost twice the specific energy during shallow cutting but generated an equal or slightly lower amount of primary dust than did the reference bit. Incendivity was eliminated for two radial designs. One conical design had ignition only after more than 15 impacts, but the other conical design was incendive. Impact failure occurred, but not so quickly as anticipated. Over 7,500 impacts on a sandstone face, with a total cutting distance of more than 1,981.2 m, only scuffed the leading edge of the sintered diamond cutting face but put hairline cracks in the substrate mounting pad. Additional design and testing are needed to develop an optimum cutter, but prospects look good for this material.

Jan 1, 1983

By Zachary Q. Maassen

The purpose of this paper is to highlight the environmental benefits of properly managing slurry waste as it is generated on site. The following projects will be used to illustrate the importance of proper onsite waste management. Project: Jet Grout Backflow Dewatering The goal was to dewater the backflow directly onsite to produce a dry, solid waste and reusable water. The backflow spoil was pumped directly to the dewatering system as it was generated. Utilizing mechanical separation equipment and flocculation polymer the waste volume was reduced by 44%. Due to the site footprint and high volumes, slurry waste removal and onsite processing were the only two feasible options. Onsite dewatering proved to be the environmentally beneficial option as it eliminated the need for slurry waste hauling and disposal. The benefits of utilizing a dewatering system are listed below: Project: Soil Mixing Spoil Recycling for Slurry Wall Construction In this application the spoils generated contain bentonite mud, cement mixture and organic ground soils. The goal of this project was to process the backflow mixture to recover reusable bentonite mud and separate the soil as a dry, stackable waste product. The treatment system replaced the traditional method of vacuum truck extraction and facility disposal. The environmental benefits of properly handling the backflow is a decrease in fresh water usage, bentonite additive reduction, and the reduction of CO2 emissions. The paper will consist of a project scope, technical equipment and process description, environmental impact reduction calculations and lessons learned for each project.

Jan 1, 2018

By Serge Ouellet

Identification and quantification of cementitious phases in cemented paste backfill (CPB) is difficult due to the fineness of tailings and the low binder content used in the mixture. Techniques such as X-ray diffraction and thermogravimetry were ineffective (mainly due to tailing mineral interference) to identify and quantify hydration products in most works reported in the literature. On the other hand, scanning electron microscopy (SEM) was used in many cases but mainly for qualitative purposes. This paper presents the development of a new approach based on a computer program (SEM-XMAP) that combines X-ray dot mapping images acquired by SEM and generates a coloured image representing the mineralogy and the microstructure of a given CPB specimen. The technique is applied to one CPB sample mixed with blast furnace slag (BFS) and acidic sulphated mixing water sampled in a mine dewatering system. The technique allowed identifying calcic matrix, unhydrated and partially hydrated BFS particles, and sulphated phases.

May 1, 2007

By A. S. Moghaddam

The transportation of cemented and non-cemented mine tailings in the form of paste through pipelines is one of the main stages of paste backfill operations. For the purpose of pipeline design, the rheological properties of the materials, such as yield stress and viscosity must be determined. In this study, the yield stress of two Canadian mine tailings, for gold and copper mineralization, were studied using the vane method and slump tests. The comparison between the vane method and the slump tests indicates that the slump tests generally overestimate the yield stress values for both cemented and non-cemented paste backfills. In addition, the effects of 5% Portland cement (type 10) and mixing time on the yield stress of paste backfills were investigated. The percentage of cement as a binder agent has a direct effect on the increase of the yield stress of paste backfills. Yield stress measurements show that mixing times of 4 minutes and 12 minutes have the same effect on the yield stress values.

May 1, 2007

By T. Belem

The main objective of this study is to better understand the effect of underground CPB placement conditions on its microstructural, hydro-mechanical and geochemical properties. To reach this objective, three PVC columns, each 3 m high, were built, and were filled with CPB material. The columns allow simulating undrained (UD), half-drained (HD) and fully-drained (FD) conditions along with the measurement of resulting self-weight consolidation settlement of the CPB. Bleed and drainage waters were also collected from each column for geochemical analyses (pH, Eh, EC, SO42- content). After 90-days of curing time, a total of sixty CPB samples were cored from the backfill columns and subjected to uniaxial compression tests followed by physical characterization (Gs, void ratio, water content, degree of saturation, specific surface area). The results showed that the total drainage water and the maximum observed settlement occur mainly within the first 72 hours after the columns are filled. The corresponding volumetric strains suggest that in situ backfilled stopes simulated behave in a similar way to the fully-drained (FD) or the half-drained (HD) conditions. The MIP test results show that the pore structure varies between the top and the bottom of each column.

May 1, 2007

By D. J. Hallbom

This paper presents a novel methodology for the hydraulic design of mine backfill distribution systems. The ?operating range? method uses the concepts of the hydraulic grade line and equivalent pipe length to determine flows and pressures as the pipeline route and static head change over the life of the mine. The operating range method is easy to learn because it is semi-graphical and uses well-known design concepts (e.g., the hydraulic grade line) and slurry design models (e.g., the Durand deposition velocity model). However, by changing the paradigm in which these concepts are used, the operating method allows all potential pipeline cases to be analysed simultaneously without the need to know the exact pipe routing. It also allows the hydraulics engineer to quickly identify ?pinch points? and ?slack flow? zones, to estimate the effect of changing the size or length of any given section of pipe, to assess the impact of changing the material properties of the backfill slurry, to size the slurry pumps, and to size and locate rupture discs when required.

May 1, 2007

By D. W. Scott

Ontario has a diverse industrial mineral resource base including abundant resources of structural industrial minerals such as sand and gravel, crushed stone, building stone, clay, shale, limestone and dolostone. These resources are used in the production of aggregates, brick, metallurgical stone, fillers, lime, cement and other construction materials. The majority of the operations are in deposits of Paleozoic and Quaternary age in southern Ontario, close to the major markets. Province-wide, there were over 140 million metric tonnes of aggregates produced in 1985, approximately 115 million tonnes of which was produced in southern Ontario with the remainder being largely from Crown land in northern Ontario. The structural industrial minerals account for over 80% of the value of total industrial mineral production, with the remaining 20% of the value king accounted for by the non-structural minerals being led by salt which is produced in southwestern Ontario. A host of other non-structural industrial minerals are known but production is currently restricted to calcite, gypsum, quartz, talc, nepheline syenite and by-product metallurgical sulphur products (Fig. 1). Location and Market Area Major production is in the area closest to the large markets of urbanized southern Ontario and the northeastern and north-central United States. In general, with the abundant variety and widespread resources of industrial minerals and the large population within 400 miles of most deposits, the market offers great potential. Ontario's 9 million population accounts for about 35% of the entire gross domestic product of Canada of $476.361 billion dollars in 1985. This equates to about 113 of the nation's buying power being within a 200 km (125 miles) radius of Toronto. This major concentration of industrial and economic activity has resulted in a well developed industrial minerals industry fulfilling most of the structural industrial minerals demand in Ontario, as well as a subsidiary demand in neighbouring provinces and states.

Jan 1, 1987

By Minerals and Materials Analysis Section

During 1996, US mines produced 319 t (10.26 million or) of gold metal valued at $4 billion. At yearend, mine reserves were estimated to be 4.7 kt (151 million oz) of gold metal. Almost all of the production was from western states. Nevada accounted for about 68% of national production. Information for this article was obtained through annual reports of mining companies, US Geological Survey (USGS) databases and other public sources. Table 1 shows production and reserves by mine. It divides the mines into two categories. These are producing mines whose primary product is gold, and producing mines for which gold is a byproduct or coproduct. The mines listed in Table 1 accounted for more than 98% of US gold production during 1996. Except the Daisy Mine, the table does not include mines that produced less than 31 kg/a (1,000 oz/year). The Daisy Mine started production in December of 1996. Nor does it contain mines whose byproduct concentrate is

Jan 1, 1997